The simplest grinding circuit consists of a ball or rod mill in closed circuit with a classifier; the flow sheet is shown in Fig. 25 and the actual layout in Fig. 9. This single-stage circuit is chiefly employed for coarse grinding when a product finer than 65 mesh is not required, but it can be adapted for fine grinding by substituting a bowl classifier for one of the straight type so as to enable the W/S ratio of the overflow to be kept below the 4/1 limit usually necessary for flotation. On account of the greater efficiency of the bowl classifier the trend of practice is towards its installation in plants grinding as coarse as 65 mesh.
Single-stage grinding is generally to be recommended for small plants on account of its simplicity. Variations in the size and character of the ore are unavoidable in most plants, but they are, as a rule, very much more noticeable when operations are on a small than when they are on a large scale. Multi-stage grinding as practised in large installations may, therefore, prove impossible to control on a very small scale, and for this reason the simplicity of single-stage grinding is likely to result in a greater overall efficiency than would be obtained with a multi-stage arrangement. When, however, the capacity of a plant approaches or exceeds 1,000 tons per day, two-stage grinding becomes preferable because the effect of normal variations of the ore is less marked and control becomes correspondingly easier.
The usual type of two-stage circuit is shown in Fig. 26, and is one that can be employed for any degree of grinding, although a straight must be substituted for a bowl classifier in the second stage when a 48-mesh product is required. It used to be the practice at one time to omit the first classifier and to pass the feed straight through the primary mill to the secondary circuit, but it was not a good method because either the secondary mill received pieces of ore that were too big or else the primary mill overground a large proportion of the feed. Much better results are obtainable by keeping the coarse ore circulating round the primary circuit, which is set for the efficient grinding of such material, until it is fine enough to be sent to the secondary circuit where the machines are set to grind fine ore more efficiently than coarse. It should be noted that the overflow of the primary classifier is sent to the secondary classifier, not direct to the mill, in order that all material which has been ground fine enough in the primary circuit can be discharged immediately without any chance of its entering and being overground in the secondary ball mill.
So important is it from the point of view of efficiency to get the undersize out of the circuit at the earliest possible moment, whether it is produced in the primary or in the secondary mill, that a special intermediate bowl classifier is often installed between the two stages. Such an arrangement has been found very useful in plants in which improvements in dry crushing practice have resulted in a reduction in the size of the feed to the grinding mills with the result that they have been able to take larger tonnages; the classifiers have then becomeoverloaded, especially in the case of older installations in which both stages were equipped with straight classifiers.
The method of installing a bowl classifier to overcome the difficulty is shown in Fig. 27. This circuit is usually adopted in modern practice, but with a bowl instead of a straight classifier, if necessary, in the closed circuit of the secondary ball mill.
Any increase in the efficiency of classification gives greater economy of power by reducing the amount of ore that is overground, so making a larger proportion of the power required to turn the mill available forgrinding the particles that are still too large. From a theoretical point of view, the ideal method of grinding would consist of a series of ball mills, each in closed circuit with a classifier and each so short that the ore in its passage through the mill would be struck only two or three times by the balls, any undersize produced being removed at once by the classifier ; in this way the chance of a particle being struck again after it had reached the required size would be reduced to a minimum. Practice circuit design approaches the ideal:
Open circuit grinding consists of one or more grinding mills, either parallel or in series, that discharges a finalground product without classification equipment and no return of coarse discharge back to the mill. Some very simplistic examples of open circuit grinding are see below and are made of aRod mill,Ball Mill or aRod mill, ball mill combination.
Not all ores can be ground in an open circuit type ofarrangement. Some conditions which do favor open circuit grinding such assmall reduction ratios,reduction of particles to a coarse, natural grainsize,recirculation of cleaner flotation middlings forregrinding anda non-critical size distribution of the final groundproduct.
Closed circuit grinding consists of one or more mills discharging ground product to classifiers which in turn return the coarse product from the size separation back to the mill for further grinding. In this circuit, grinding efficiency is very dependent upon the size separation effected so care should be exercised in selecting the type and size of classifier used to close the system.
This type of grinding is the most common circuit found in mineral processing facilities, mainly because a lot of ores and product requirements are not suitable for open circuit grinding. Some advantages presented by grinding in closed circuit are that this arrangement usually results in higher mill capacity and lower power consumption per ton of product, it eliminates overgrinding by removing fines early and it avoids coarse material in the final ground product by returning this material to the mill.
Although closed circuit grinding offers many choices for arrangement of the equipment as well as combinations of equipment, some of the more common circuits arerod mill/classifier,Ball mill/Classifier,Rod mill/Ball mill/Classifier and Rod mill/Classifier/Ball mill/Classifier.
The importance of the grinding circuit to overall production in any facility should be obvious by now. Because of the responsibilities assigned to grinding it becomes essential that a grinding mill accepts a certain required tonnage of ore per day while yielding a product that is of a known and controllable particle size. This leads to the conclusion that close control over the grinding circuit is extremely important.
There are many factors which can contribute to fluctuationsin performance of a mill, but some of the most common found inindustrial practice are thechanges in ore taken from different parts of the mine,changes in crusher settings,wear in the crushers,screen damage in the crusher circuit.
These are a few things that operators should look for when changes in mill performance are noticed. Stockpiling of ore ahead of the mill can aid in smoothing out some of the fluctuations although it must be stored in such a manner that no segregation occurs.
The reduction ratio in the grinding section is so much greater than in the crushing plant that labour becomes a relatively small item and the power and steel consumption the largest items of cost. Table 18 gives the average total consumption of power that may be expected in modern ball mill installations of various capacities up to 4,000 tons per day, the figures being based on an average medium-hard ore.
The cost of grinding is more difficult to predict than that of crushing because variations in the hardness and toughness of the ore produce proportionately wider variations in the consumption of power and steel. An approximate guide to grinding costs; they are direct costs and include no overhead charges. Power is assumed to cost 0.075 per kilowatt-hour in the case of the smallest plant and to decrease to a minimum per kilowatt-hour for the largest.
As the tonnage rises up to 1,000 tons per day the costs fall rapidly. In plants of greater capacity, however, they do not decrease in the same proportion with increase of tonnage, because the extra capacity is not obtained by increasing the size of the individual machines but by installing two or more similar units side by side, each of equal efficiency.Reduction of costs then becomes more a matter of organization than of plant design.
As already stated, the power and steel costs are the two largest items, those of labour and supplies being small by comparison ; it is on this account that recent progress has been mainly directed towards reducing the consumption of power and steel by means of greater efficiency in classification and by the use of mills of larger diameter.
The way in which the efficiency of classification has been increased has been described in the paragraph headed Grinding Circuits. An increase in the diameter of a mill gives greater economy in two ways : In the first place, the balls do more effective work in a large than in asmall mill, because, falling from a greater height, they shatter the pieces of ore with greater force ; in the second place, the ratio of the deadweight of the mill to the weight of the ball charge decreases as the diameter increases and thus in a large mill the useless weight to be moved is distributed over a greater weight of useful ball load than in a small mill, with the result that a larger proportion of the total power consumption is available to give the balls the cascading and rolling action necessary to break up the ore.
It is essential for the grinding and flotation sections of a plant to be run continuously. It takes nearly half an hour to clear the circuit of even a small grinding unit preparatory to stopping it, and often an hour is necessary to get the circuit fully loaded after restarting ; most of the power required to keep the machines running during the stopping and starting periods is wasted. Moreover, the operation of the flotation machines is poor during these periods so that much of the power required to keep them running is also wasted. In addition, modern practice aims at the elimination of everything likely to cause fluctuating conditions. For these reasons it is the universal custom to run the grinding and flotation plants for 24 hours per day.
Grinding circuits are fed at a controlled rate from the stockpile or bins holding the crusher plant product. There may be a number of grinding circuits in parallel, each circuit taking a definite fraction of the feed. An example is the Highland Valley Cu/Mo plant with five parallel grinding lines (Chapter 12). Parallel mill circuits increase circuit flexibility, since individual units can be shut down or the feed rate can be changed, with a manageable effect on production. Fewer mills are, however, easier to control and capital and installation costs are lower, so the number of mills must be decided at the design stage.
The high unit capacity SAG mill/ball mill circuit is dominant today and has contributed toward substantial savings in capital and operating costs, which has in turn made many low-grade, high-tonnage operations such as copper and gold ores feasible. Future circuits may see increasing use of high pressure grinding rolls (Rosas et al., 2012).
Autogenous grinding or semi-autogenous grinding mills can be operated in open or closed circuit. However, even in open circuit, a coarse classifier such as a trommel attached to the mill, or a vibrating screen can be used. The oversize material is recycled either externally or internally. In internal recycling, the coarse material is conveyed by a reverse spiral or water jet back down the center of the trommel into the mill. External recycling can be continuous, achieved by conveyor belt, or is batch where the material is stockpiled and periodically fed back into the mill by front-end loader.
In Figure 7.35 shows the SAG mill closed with a crusher (recycle or pebble crusher). In SAG mill operation, the grinding rate passes through a minimum at a critical size (Chapter 5), which represents material too large to be broken by the steel grinding media, but has a low self-breakage rate. If the critical size material, typically 2550mm, is accumulated the mill energy efficiency will deteriorate, and the mill feed rate decreases. As a solution, additional large holes, or pebble ports (e.g., 40100mm), are cut into the mill grate, allowing coarse material to exit the mill. The crusher in closed circuit is then used to reduce the size of the critical size material and return it to the mill. As the pebble ports also allow steel balls to exit, a steel removal system (such as a guard magnet, Chapters 2 and 13Chapter 2Chapter 13) must be installed to prevent them from entering the crusher. (Because of this requirement, closing a SAG mill with a crusher is not used in magnetic iron ore grinding circuits.) This circuit configuration is common as it usually produces a significant increase in throughput and energy efficiency due to the removal of the critical size material.
An example SABC-A circuit is the Cadia Hill Gold Mine, New South Wales, Australia (Dunne et al., 2001). The project economics study indicated a single grinding line. The circuit comprises a SAG mill, 12m diameter by 6.1m length (belly inside liners, the effective grinding volume), two pebble crushers, and two ball mills in parallel closed with cyclones. The SAG mill is fitted with a 20MW gearless drive motor with bi-directional rotational capacity. (Reversing direction evens out wear on liners with symmetrical profile and prolongs operating time.) The SAG mill was designed to treat 2,065t h1 of ore at a ball charge of 8% volume, total filling of 25% volume, and an operating mill speed of 74% of critical. The mill is fitted with 80mm grates with total grate open area of 7.66m2 (Hart et al., 2001). A 4.5m diameter by 5.2m long trommel screens the discharge product at a cut size of ca. 12mm. Material less than 12mm falls into a cyclone feed sump, where it is combined with discharge from the ball mills. Oversize pebbles from the trommel are conveyed to a surge bin of 735t capacity, adjacent to the pebble crushers. Two cone crushers with a closed side set of 1216mm are used to crush the pebbles with a designed product P80 of 12mm and an expected total recycle pebble rate of 725t h1. The crushed pebbles fall directly onto the SAG mill feed belt and return to the SAG mill.
SAG mill product feeds two parallel ball mills of 6.6m11.1m (internal diameterlength), each with a 9.7MW twin pinion drive. The ball mills are operated at a ball charge volume of 3032% and 78.5% critical speed. The SAG mill trommel undersize is combined with the ball mills discharge and pumped to two parallel packs (clusters) of twelve 660mm diameter cyclones. The cyclone underflow from each line reports to a ball mill, while the cyclone overflow is directed to the flotation circuit. The designed ball milling circuit product is 80% passing 150m.
Several large tonnage copper porphyry plants in Chile use an open-circuit SAG configuration where the pebble crusher product is directed to the ball mills (SABC-B circuit). The original grinding circuit at Los Bronces is an example: the pebbles generated in the two SAG mills are crushed in a satellite pebble crushing plant, and then are conveyed to the three ball mills (Mogla and Grunwald, 2008).
Hydrocyclones have come to dominate classification when dealing with fine particle sizes in closed grinding circuits (<200m). However, recent developments in screen technology (Chapter 8) have renewed interest in using screens in grinding circuits. Screens separate on the basis of size and are not directly influenced by the density spread in the feed minerals. This can be an advantage. Screens also do not have a bypass fraction, and as Example 9.2 has shown, bypass can be quite large (over 30% in that case). Figure 9.8 shows an example of the difference in partition curve for cyclones and screens. The data is from the El Brocal concentrator in Peru with evaluations before and after the hydrocyclones were replaced with a Derrick Stack Sizer (see Chapter 8) in the grinding circuit (Dndar et al., 2014). Consistent with expectation, compared to the cyclone the screen had a sharper separation (slope of curve is higher) and little bypass. An increase in grinding circuit capacity was reported due to higher breakage rates after implementing the screen. This was attributed to the elimination of the bypass, reducing the amount of fine material sent back to the grinding mills which tends to cushion particleparticle impacts.
Circulation of material occurs in several parts of a mineral processing flowsheet, in grinding and flotation circuits, for example, as well as the crushing stage. In the present context, the circulating load (C) is the mass of coarse material returned from the screen to the crusher relative to the circuit final product (or fresh feed to the circuit), often quoted as a percentage. Figure 8.2 shows two closed circuit arrangements. Circuit (a) was considered in Chapter 6 (Example 6.1), and circuit (b) is an alternative. The symbols have the same meaning as before. The relationship of circulating load to screen efficiency for circuit (a) was derived in Example 6.1, namely (where all factors are as fractions):
The circulating load as a function of screen efficiency for the two circuits is shown in Figure 8.3. The circulating load increases with decreasing screen efficiency and as crusher product coarsens (f or r decreases), which is related to the crusher set (specifically the closed side setting, c.s.s.). For circuit (a) C also increases as the fresh feed coarsens (n decreases), which is likely coming from another crusher. In this manner, the circulating load can be related to crusher settings.
In industrial grinding process, in addition to goal of productivity maximization, other purposes of deterministic grinding circuit optimization have to satisfy the upper bound constraints on the control variables. We know that there lies a tradeoff between the throughput (TP) and the percent passing of midsize classes (MS) from the previous work of Mitra and Gopinath,2004. In deterministic optimization formulation, there are certain parameters which we will assume them as constant. But, in real life that may not be case. There are such six parameters in our industrial grinding process which are R, B, R, B are the grindability indices and grindability exponents for the rod mill (RMGI) and the ball mill (BMGI); and P, S are the sharpness indices for the primary (PCSI) and secondary cyclones (SCSI). These parameters are treated as constant in deterministic formulation. As they are going to be treated as uncertain parameters in the OUU formulation. These parameters are assumed uncertain because most of them are obtained from the regression of experimental data and thus are subject to uncertainty due to experimental and regression errors. In the next part of the section, we consider them as fuzzy numbers and solve the OUU problem by FEVM. In FEVM formulation, the uncertain parameters are considered as fuzzy numbers and the uncertain formulation is transformed into the deterministic formulation by expectation calculations for both objective function and constraints. So, the converted deterministic multi-objective optimization problem is expressed as:
Another spinning batch concentrator (Figure 10.27), it is designed principally for the recovery of free gold in grinding circuit classifier underflows where, again, a very small (<1%) mass pull to concentrate is required. The feed first flows up the sides of a cone-shaped bowl, where it stratifies according to particle density before passing over a concentrate bed fluidized from behind by back-pressure (process) water. The bed retains dense particles such as gold, and lighter gangue particles are washed over the top. Periodically the feed is stopped, the bed rinsed to remove any remaining lights and is then flushed out as the heavy product. Rinsing/flushing frequency, which is under automatic control, is determined from grade and recovery requirements.
The units come in several designs, the Semi-Batch (SB), Ultrafine (UF), and i-Con, designed for small scale and artisanal miners. The first installation was at the Blackdome Gold Mine, British Columbia, Canada, in 1986 (Nesset, 2011).
These two batch centrifugal concentrators have been widely applied in the recovery of gold, platinum, silver, mercury, and native copper; continuous versions are also operational, the Knelson Continuous Variable Discharge (CVD) and the Falcon Continuous (C) (Klein et al., 2010; Nesset, 2011).
To liberate minerals from sparsely distributed and depleting the ore bodies finer grinding than generally obtained by the conventional Rod Mill Ball Mill grinding circuits is needed. Longer grinding periods in the conventional milling processes prove too expensive mainly due to large power consumption. Stirrer mills have been tried in mineral industry with considerable success and have therefore been increasingly used. In this chapter, the theories involved in the design and operation of these mills, as established till now, are explained. Further theoretical studies and designs of the mills are still in progress for a better understanding and improved operation. Presently, the mills have been proved to be economically viable and the mineral of interest conducive to improved recovery and grade.
IMP Technologies Pty. Ltd. has recently tested a pilot-scale super fine crusher that operates on dry ore and is envisaged as a possible alternative to fine or ultra-fine grinding circuits (Kelsey and Kelly, 2014). The unit includes a rotating compression chamber and an internal gyrating mandrel (Figure 6.13). Material is fed into the compression chamber and builds until the gyratory motion of the mandrel is engaged. Axial displacement of the compression chamber and the gyratory motion of the mandrel result in fine grinding of the feed material. In one example, a feed F80 of 300m was reduced to P80 of 8m, estimated to be the equivalent to two stages of grinding. This development is the latest in a resurgence in crushing technology resulting from the competition of AG/SAG milling and the demands for increased comminution energy efficiency.
The iron oxide crystal grains in most iron ores are not evenly distributed in size. Spiral separators can therefore be used to take out the coarser iron oxide grains in the primary grinding circuit to save grinding energy and help achieve a higher iron recovery. Figure 9.14 presents a typical flow sheet for processing an oxidized ore containing about 30% Fe using a combination of spiral and SLon magnetite separators and reverse flotation. This ore is mainly composed of hematite, magnetite, and quartz, and the iron oxide crystals range in size from 0.005 to 1.0mm with an average size of about 0.05mm. The average size of the quartz crystals is approximately 0.085mm.
In the primary grinding stage of the flow sheet in Figure 9.14, the ore is first ground down to about 60% -75m and then classified into two size fractions, a coarse size fraction and a fine size fraction. The coarse size fraction is treated with spiral separators to recover part of the final iron ore concentrate. Then, drum LIMS and SLon magnetic separators are used to reject some of the coarse gangue minerals as final tailings. The magnetic products from the LIMS and SLon are sent back to the secondary ball mill for regrinding, and the milled product returns to the primary cyclone classifier.
The fine size fraction is about 90% -75m and is processed using drum LIMS separators and SLon magnetic separators in series to take out the magnetite and hematite, respectively. The magnetic products from the magnetic separators are mixed to generate the feed for reverse flotation to produce another component of the final iron ore concentrate.
The key advantage of this flow sheet lies in the fact that the spirals and SLon magnetic separators take out about 20% of the mass of the final iron concentrate and about 20% of the mass of the final tailings, respectively, from the coarse size fraction. This greatly reduces the masses being fed to the secondary ball mill and reverse flotation, thereby greatly reducing the total processing cost. From the plant results for this flow sheet, an iron concentrate containing 67.5% Fe could be produced from a run-of-mine ore containing 30.1% Fe, at a mass yield to the iron concentrate of 34.9%, an iron recovery of 78.0%, and a tailings grade of 10.2% Fe.
The first step of physical beneficiation is crushing and grinding the iron ore to its liberation size, the maximum size where individual particles of gangue are separated from the iron minerals. A flow sheet of a typical iron ore crushing and grinding circuit is shown in Figure 1.2.2 (based on Ref. ). This type of flow sheet is usually followed when the crude ore contains below 30% iron. The number of steps involved in crushing and grinding depends on various factors such as the hardness of the ore and the level of impurities present .
Jaw and gyratory crushers are used for initial size reduction to convert big rocks into small stones. This is generally followed by a cone crusher. A combination of rod mill and ball mills are then used if the ore must be ground below 325 mesh (45m). Instead of grinding the ore dry, slurry is used as feed for rod or ball mills, to avoid dusting. Oversize and undersize materials are separated using a screen; oversize material goes back for further grinding.
Typically, silica is the main gangue mineral that needs to be separated. Iron ore with high-silica content (more than about 2%) is not considered an acceptable feed for most DR processes. This is due to limitations not in the DR process itself, but the usual customer, an EAF steelmaking shop. EAFs are not designed to handle the large amounts of slag that result from using low-grade iron ores, which makes the BF a better choice in this situation. Besides silica, phosphorus, sulfur, and manganese are other impurities that are not desirable in the product and are removed from the crude ore, if economically and technically feasible.
While used sometimes on final concentrates, such as Fe concentrates, to determine the Blaine number (average particle size deduced from surface area), and on tailings for control of paste thickeners, for example, the prime application is on cyclone overflow for grinding circuit control (Kongas and Saloheimo, 2009). Control of the grinding circuit to produce the target particle size distribution for flotation (or other mineral separation process) at target throughput maximizes efficient use of the installed power.
Continuous measurement of particle size in slurries has been available since 1971, the PSM (particle size monitor) system produced then by Armco Autometrics (subsequently by Svedala and now by Thermo Gamma-Metrics) having been installed in a number of mineral processing plants (Hathaway and Guthnals, 1976).
The PSM system uses ultrasound to determine particle size. This system consists of three sections: the air eliminator, the sensor section, and the electronics section. The air eliminator draws a sample from the process stream and removes entrained air bubbles (which otherwise act as particles in the measurement). The de-aerated pulp then passes between the sensors. Measurement depends on the varying absorption of ultrasonic waves in suspensions of different particle sizes. Since solids concentration also affects the absorption, two pairs of transmitters and receivers, operating at different frequencies, are employed to measure particle size and solids concentration of the pulp, the processing of this information being performed by the electronics. The Thermo GammaMetrics PSM-400MPX (Figure 4.18) handles slurries up to 60% w/w solids and outputs five size fractions simultaneously.
Other measurement principles are now in commercial form for slurries. Direct mechanical measurement of particle size between a moving and fixed ceramic tip, and laser diffraction systems are described by Kongas and Saloheimo (2009). Two recent additions are the CYCLONEtrac systems from CiDRA Minerals Processing (Maron et al., 2014), and the OPUS ultrasonic extinction system from Sympatec (Smith et al., 2010).
CiDRAs CYCLONEtrac PST (particle size tracking) system comprises a hardened probe that penetrates into the cyclone overflow pipe to contact the stream and effectively listens to the impacts of individual particles. The output is % above (or below) a given size and has been shown to compare well with sieve sizing (Maron et al., 2014). The OPUS ultrasonic extinction system (USE) transmits ultrasonic waves through a slurry that interact with the suspended particles. The detected signal is converted into a particle size distribution, the number of frequencies used giving the number of size classes measured. Applications on ores can cover a size range from 1 to 1,000m (Smith et al., 2010).
In addition to particles size, recent developments have included sensors to detect malfunctioning cyclones. Westendorf et al. (2015) describe the use of sensors (from Portage Technologies) on cyclone overflow and underflow piping. CiDRAs CYCLONEtrac OSM (oversize monitor) is attached to the outside of the cyclone overflow pipe and detects the acoustic signal as oversize particles (rocks) hit the pipe (Cirulis and Russell, 2011). The systems are readily installed on individual cyclones thus permitting poorly operating units to be identified and changed while allowing the cyclone battery to remain in operation. Figure 4.19 shows an installation of both CiDRA systems (PST, OSM) on the overflow pipe from a cyclone.